TN295 
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No. 9176 



LIBRARY OF CONGRESS 



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Bureau of Mines Information Circular/1988 




Processing Technologies for Extracting 
Cobalt From Domestic Resources 



By C. E. Jordan 




UNITED STATES DEPARTMENT OF THE INTERIOR 




Information Circular 9176 



Processing Technologies for Extracting 
Cobalt From Domestic Resources 



By C. E. Jordan 



UNITED STATES DEPARTMENT OF THE INTERIOR 
Donald Paul Hodel, Secretary 

BUREAU OF MINES 

David S. Brown, Acting Director 



TV 



Library of Congress Cataloging in Publication Data: 



Jordan, C. E. 

Processing technologies for extracting cobalt from domestic resources. 



(Bureau of Mines information circular; 9176) 

Bibliography: p. 23-24. 

Supt. of Docs, no.: I 28.27:9176. 

1. Cobalt— Metallurgy. 2. Cobalt ores— United States. I. Title. II. Series: Information 
circular (United States. Bureau of Mines) ; 9176. 

TN296.U4 [TN799.C6] 622 s [669'.733] 87-600331 



CONTENTS 



Page 



Abstract 

Introduction 

Sulfide ores 

Blackbird deposit 

Duluth Gabbro deposits 

Yakobi Island deposit 

Madison Mine 

Missouri lead belt deposits 

Copper deposits 

Pyrite concentrates 

Spent copper leach solutions 

Iron deposits 

Oxides 

Laterite deposits 

Manganese sea nodules 

Extraction methods. 

Gas reduction and ammoniacal leach process 

Cuprion ammoniacal leach process 

High -temperature, high-pressure H 2 S0 4 leach process 

Reduction and HC1 leach process 

Smelting and H 2 S0 4 leach process 

Refining process 

Ammoniacal solution refining 

Acid sulfate refining 

Acid chloride refining 

Manganese sea crusts 

Discussion 

Conclusions 

References 

ILLUSTRATIONS 

1. Domestic cobalt resources 3 

2. Blackbird Mine cobalt process 5 

3. Madison Mine cobalt process 8 

4. Cobalt from spent Cu leach solution 10 

5. Cobalt from Pennsylvania iron ore 11 

6. Reduction roast -ammoniacal leach process for laterites 13 

7. Sulfuric acid leach process for laterites 14 

8. Gas reduction and ammoniacal leach process for sea nodules 16 

9. Cuprion ammoniacal leach process for sea nodules 16 

10. High-temperature, high-pressure H2SO4 leach process for sea nodules 17 

11. Reduction and HC1 leach process for sea nodules 17 

12. Smelting and H2SO4 leach process for sea nodules 18 

TABLE 
1. U.S. cobalt resource summary 22 



1 


2 


4 


4 


6 


7 


7 


7 


9 


9 


9 


11 


12 


12 


15 


15 


15 


15 


15 


15 


18 


18 


18 


19 


19 


20 


20 


23 


23 





UNIT OF MEASURE 


ABBREVIATIONS 


USED 


IN THIS 


REPORT 


cm 


centimeter 






pet 


percent 


°C 


degree Celsius 






ppm 


part per million 


g/L 


gram per liter 






psi 


pound per square inch 


h 


hour 






st/d 


short ton per day 


lb 


pound 






um 


micrometer 


lb/h 


pound per hour 






yr 


year 


mm Hg 


millimeter of mercury 









PROCESSING TECHNOLOGIES FOR EXTRACTING 
COBALT FROM DOMESTIC RESOURCES 

by C. E. Jordan 1 



ABSTRACT 

Domestic cobalt resources are relatively large, but low grade. The 
full potential for cobalt production from domestic sources is at least 
19.4 million lb of cobalt per year exclusive of offshore resources. A 
summary of the cobalt processing technologies for the major domestic 
resources is presented in this Bureau of Mines report. The processing 
technologies for the Blackbird, Madison Mine, Duluth Gabbro, iron ore 
pyrite, laterites, and manganese sea nodules are nearly complete, but 
the economics are not favorable. Research on these resources should be 
limited to approaches that promise to cut the total processing costs by 
at least 50 pet. The most promising sources of cobalt are the spent 
copper leach solutions and siegenite from the Missouri lead ores. 
Research on cobalt processing from these two sources needs to be 
completed. 



'Metallurgist, Tuscaloosa Research Center, Tuscaloosa, AL. 



INTRODUCTION 



The United States is the largest 
consumer of cobalt in the world, but has 
no domestic cobalt production. Except 
for some scrap recycling, the United 
States depends entirely on foreign 
nations to supply over 15 million lb of 
cobalt annually. More than half of the 
cobalt metal comes from Zaire and Zambia 
(19). 2 With most of the cobalt coming 
from a small number of foreign producers, 
the United States is vulnerable to supply 
disruptions. 

Cobalt is a critical element in many 
industrial and military products such as 
jet engine parts, high-strength tool 
steels, heat- and corrosion-resistant 
alloys, magnets, catalysts, drying addi- 
tives in paints, and other chemicals. 
About two-thirds of the domestic cobalt 
consumption requires cobalt metal, either 
as a powder or as a high-purity cathode. 
The remaining third is as oxides or salts 
used in chemicals and paint drying addi- 
tives (19). About half of the 15 million 
lb the United States consumes annually is 
considered essential (22). In recogni- 
tion of the strategic importance of 
cobalt, research has been directed in 
three areas, recycling technology, cobalt 

— - 

Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this report. 



substitutes, and recovery technology from 
domestic resources. Currently, recycling 
accounts for about 8 pet of domestic 
consumption and has a potential for over 
25 pet (19). Substitution technology has 
potential for an additional 25 pet, but 
research and development is moving slowly 
owing to the current low price of cobalt 
(22). Extensive research on primary 
cobalt production has been conducted on 
virtually every major cobalt-bearing 
deposit. In an effort to optimize future 
research and establish research priori- 
ties, the Bureau of Mines investigated 
the status of processing technologies for 
extracting cobalt from primary domestic 
resources. 

Because domestic cobalt resources are 
low grade, cobalt production is likely 
only if metals such as nickel, copper, 
iron, lead, and zinc are recovered. 
There are two basic types of cobalt 
deposits, sulfide and oxide. The benefi- 
ciation and processing largely depends 
upon the type of ore and the associated 
metals. Each resource with over 15 
million lb of Co is presented in this 
report (fig. 1). Along with the resource 
description, the benef iciation and 
processing technologies are presented 
with an assessment of the technical 
progress and the environmental problems. 



CA 



OR 



ID 



NV 



MT 



WY 



ND MINI 



SD 



NE 



IA 



Wl 



IL 



CO 








KS *> 






OK 


NM 




rx 



jyio_ 

AR 



NH- 



ME 



J// 



Ml 



IN 



OH 



PA 



KY 



TN, 



MS 



LA 



AL 



,GA 



WV. 



SC 



VA 



NC 



-MA 
Rl 



NJ 
-DE 

MD 



If 


o 




\ 




\ 


Cb 


\ 
\ 






~v 


\ 

X 


sHI N 


s 



..^a:*? 1 ^? 




LEGEND 
• Cobalt 



FL 



FIGURE 1 .—Domestic cobalt resources. 



SULFIDE ORES 



BLACKBIRD DEPOSIT 

The Blackbird deposit near Salmon, ID, 
contains 0.65 pet Co with nearly all of 
the cobalt found in cobaltite (CoAsS) and 
a minor amount associated with chalcopy- 
rite (CuFeS 2 ). The deposit's proven 
reserves are over 60 million lb of Co. 
Noranda Exploration Inc., Cobalt, ID is 
waiting for favorable economic conditions 
to build a benef iciation and cobalt 
processing plant with a planned capacity 
of 4 million lb of Co per year. The 
arsenic content of the ore has focussed 
some attention on the need for safe min- 
ing and tailings disposal. High cobalt 
and copper levels were found in streams 
draining from the deposit's historical 
mining site. An ion exchange process was 
successfully field tested to remove both 
the cobalt and copper from these runoff 
streams. The environmental costs associ- 
ated with this deposit have been 
estimated at $3.00/lb of Co produced. 

Benef iciation of the Blackbird ore 
began with crushing and grinding the ore 
to 70 pet minus 200-mesh size. Using a 
sequential sulfide flotation process, 
chalcopyrite was floated first. Only 
5 pet of Co was lost in the chalcopyrite 
concentrate containing 26 pet Cu and 
0.65 pet Co. The remaining sulfides were 
floated, producing a bulk sulfide concen- 
trate containing 5 pet Co, 0.1 pet Ni , 
and 0.4 pet Cu. A pilot plant was oper- 
ated to demonstrate this process and 
80 pet of the cobalt was recovered in the 
concentrate (10). 

Historically, extraction of cobalt from 
the Blackbird concentrate began with a 
controlled oxidizing roast. However, 
because of the environmental problems 
associated with arsenic fumes, this 
technique is no longer considered appro- 
priate. Fortunately, there are three 
hydrometallurgical alternative methods 
that do not require the oxidation roast. 
Cobalt can be leached as a sulfate, 
chloride, or ammine complex. Because a 
single company owns the Blackbird 
deposit, only the process proposed by 
Noranda Exploration Inc. will be 



discussed here. The cobalt was dissolved 
by pressure leaching with sodium sulfate 
at 200° C and 150 psi oxygen pressure 
(fig. 2). Actually, the pyrite in the 
cobalt concentrate was oxidized, produc- 
ing H 2 S0 4 and Fe 2 (S0 4 ) 3 , which leached 
the cobalt from CoAsS. Sodium sulfate 
helped to suppress iron and arsenic 
dissolution in the autoclave. Ninty- 
seven percent of cobalt was extracted 
in a solution containing 30 g/L Co and 
100 g/L H 2 S0 4 . The leach slurry was 
cooled to 95° C and neutralized to pH 
1.5, precipitating jarosite and ferric 
arsenate. The pregnant solution also 
contained nickel and other impurities 
such as iron, arsenic, copper, and zinc. 
A semicontinuous pilot plant using 22-lb 
batches was operated to demonstrate the 
extraction process. Commercial equipment 
for this high-pressure oxidation-H 2 S04 
leach process is available. 

After extraction of the cobalt into 
solution, the refining process used 
largely depends upon the impurities and 
the final commercial cobalt product 
desired. Noranda chose to market a high- 
purity cathode cobalt. First, the hot 
pregnant solution was filtered from the 
leach residue, jarosite and ferric 
arsenate. At 50° C and 1.5 pH, H 2 S was 
added to precipitate the copper and 
arsenic as sulfides. The filtered solu- 
tion was oxidized and neutralized to pH 
5.0 at 75° C to precipitate the residual 
iron. Zinc was removed from the filtered 
and cooled solution by ion exchange at pH 
4.0 and ambient temperature. The solu- 
tion pH was then adjusted to 2.5 and the 
nickel was removed by ion exchange. 
Finally, the cobalt was precipitated as 
cobaltous hydroxide with a lime slurry. 
After filtering the solution, the cobalt- 
ous hydroxide was leached with spent 
cobalt electrolyte followed by cobalt 
electrowinning at 50° C. The cobalt 
metal was stripped from the cathode and 
traces of hydrogen were removed by vacuum 
degassing at 820° C and 200 mm Hg. All 
of the unit operations, except the zinc 
and nickel ion exchange steps, are estab- 
lished commercial technologies. 





^ 


Bulk con 


cent 


rate 




Wash 1 


iquor 










^ 


i 


\ 
\ 




Oxygen 


Pressure 
leach 




Ma CD %M««k 




w 


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■ I CIOII 




Lime 




' i 












\ 


r ^r 




V 




w Do<.i<J..» 


i) 




Jarosite 
precipitation 




L-S separation 






p 


w ncoiuuc 


Hydrogen 






1 




Mf — -. U 












WW CI Oil 


sulphide 




i 


r v 




y 




w D^^.^..„ 




Cu-As 
precipitation 




L-S separation 






p 


W I1C3IUUC 

(Cu-As sulphid< 


Limestone 




r i 














i 


r 






Oxygen 


Fe oxidation 

and 
precipitation 




L-S separation 


Residue 


► 


P 




'a 




i 














' 






Eluant 


Zinc 
ion-exchange 




Eluate to 

precipitation 

with lime 

Primary eluate 

Secondary eluate to 
precipitation with 
soda ash 


H 2 S0 4 


w 


P 




i 






^ 
' 




Eluant 


Nickel 
ion-exchange 


k* 


H 2 S0 4 


w 


P 


Lime 




r 1 


' 






J 










Cobalt 
precipitation 




L-S separation 


k* Filtr^tf 




P 




H 2 S0 4 


I i 












f 








Cobalt 
dissolution 




L-S separation 


Residue 




w 


P 


(Gypsum) 


w 




i 
















r 








Spent 
electrolyte 


Carbon 
column 








i 


' 






Cobalt 
electrowinning 


k» 


Evaporation 






^ 


P 








1 


r 









Co cathode 

FIGURE 2.— Blackbird Mine cobalt process. 



This refining process was also demon- 
strated in the semi continuous miniplant 
previously mentioned. The overall cobalt 
recovery from ore to cathodes was 75 pet. 
The solution purification process removed 
the Cu, Zn, Ni , Fe, and As impurities and 
the cobaltous hydroxide precipitation 
removed soluble impurities such as Na and 
Mg. The cobalt cathode purity was excel- 
lent with the exception of a high 
selenium level (10). 

The research on the Blackbird deposit 
is nearly complete. The extraction and 
refining technology has been demonstrated 
in a semicontinuous miniplant. The 
cobalt product met nearly all the 
required specifications and the environ- 
mental problems have been addressed. 
However, with estimated production costs 
at $25/lb of Co, mining and processing 
this ore for cobalt was uneconomical. 

DULUTH GABBRO DEPOSITS 

The largest domestic Cu-Ni-Co sulfide 
resource is associated with the Duluth 
Gabbro Complex in northeastern Minnesota. 
Almost half of the cobalt is in cobalt- 
bearing sulfides that are disseminated 
throughout the deposit in the form of 
pyrrhotite, pentlandite, cubanite, and 
chalcopyrite. The remaining cobalt is 
associated with olivine and plagioclase. 
Over 184 million lb of Co is found in the 
deposit, which contains 0.025 pet Co, 
0.18 pet Ni, and 0.79 pet Cu (24). Upon 
full development, the region could 
produce up to 5 million lb of Co per 
year. 

Cobalt benef iciation of the Duluth Gab- 
bro Complex ores has been limited to the 
sulfide minerals. The cobalt and nickel 
associated with oxides and silicates, 
about half of the deposit, has 
no benef iciation potential. The sulfide 
benef iciation began with crushing and 
grinding the ore to minus 200-mesh size. 
All the sulfides were floated with 
xanthate in a bulk concentrate. The 
xanthate collector was removed from the 
particle surfaces with lime. Then the 
concentrate was filtered and reground to 
minus 500-mesh size. The pentlandite and 
pyrrohotite were depressed and the 
chalcopyrite and cubanite were floated. 



The bulk sulfide flotation process and 
the differential flotation process were 
demonstrated in a 12-st/d pilot plant. 
About 40 pet of the cobalt was recovered 
in the bulk sulfide concentrate. After 
the differential float, about 8 pet of 
the cobalt was in the copper concentrate 
containing 19 pet Cu, 0.5 pet Ni , and 
0.05 pet Co. The remaining 32 pet of the 
cobalt was in the nickel-cobalt concen- 
trate containing 0.5 pet Ni and 0.5 pet 
Co (27). The benef iciation technology 
for the Duluth Gabbro deposit appears to 
be complete. Differential flotation 
recovered 32 pet of the cobalt while 
increasing the grade from 0.025 to 
0.5 pet. Environmental assessment of the 
mine tailings showed no potential 
environmental hazards. 

The extraction of cobalt from the 
Duluth Gabbro is possible as a byproduct 
from copper and nickel recovery. Using 
the bulk sulfide concentrate, a matte 
smelting technique was used to produce a 
low-iron Cu-Ni-Co matte. The concentrate 
was roasted at 720° C to lower the sulfur 
content, fluxed with Si02 and CaO, and 
smelted at 1,300° C, making a Cu-Ni-Co 
matte. Only 25 pet of the concentrate 
cobalt was recovered in the matte which 
left most of the cobalt with the iron in 
the slag (28). From ore to matte, only 
10 pet of the cobalt was recovered in the 
matte containing 61 pet Cu, 9 pet Ni , and 
only 0. 1 pet Co. This was too high in 
copper and too low in Ni and Co for the 
world's commercial Cu-Ni refineries, 
which usually treat mixed sulfides or 
mattes containing more Ni than Cu. The 
cobalt was not recovered efficiently. A 
survey of six commercial Co-Ni-Cu matte 
smelting operations showed that between 
30 to 60 pet of the cobalt was recovered 
in the matte. This indicates that matte 
smelting is not appropriate for efficient 
cobalt recovery. Alternative hydrometal- 
lurgical techniques need to be studied to 
establish cobalt recovery technology for 
the Duluth Gabbro concentrates. After- 
wards, continuous testing up to pilot 
plant scale is needed for economic 
evaluation. 

With half of the cobalt left in the 
flotation tailings, both the ore or the 
tailings may be suitable for solution 



mining or heap leaching with weak acid. 
Exploratory test work should be conducted 
to see if the cobalt and other metals can 
be efficiently leached. The technology 
for recovering cobalt from similar leach 
solutions has been developed and will be 
discussed later in this report ("Spent 
Copper Leach Solutions" section). Leach- 
ing the ore or the tailings may result in 
lower costs in recovering cobalt from 
this resource. 

YAKOBI ISLAND DEPOSIT 

Another large Cu-Ni-Co deposit is found 
on Yakobi Island in southeastern Alaska. 
A resource of approximately 14 million lb 
of Co is estimated for this deposit. 
Most of the cobalt is associated with 
sulfides. The deposit also contains 85 
million lb of Cu and 140 million lb of Ni 
(16). The average cobalt grade is 
0.04 pet. The deposit has been explored, 
but very little benef iciation and extrac- 
tion research has been conducted. The 
deposit is located in the National 
Wilderness Preservation System where 
future exploratory mining is restricted. 

MADISON MINE 

Cobalt and nickel mineralization is 
associated with lead deposits in south- 
eastern Missouri. The Madison mine 
located on the edge of the lead belt has 
a high-grade cobalt ore, making it 
uniquely different than the other lead- 
zinc deposits. At the Madison Mine up to 
21 million lb of Co is found in ore 
containing 0.17 pet Co. Siegenite, 
(Ni,Co)3S 4 , is the major cobalt mineral, 
but some cobalt is also found in chal- 
copyrite, sphalerite, galena, and miller- 
ite (32). The expected annual cobalt 
production by Anshutz Mining Co. was 
between 1 and 2 million lb (22), however, 
these plans are contingent upon more 
favorable cobalt prices. 

After crushing and grinding, bulk 
sulfide flotation was used to beneficiate 
the ore. The method was demonstrated by 
processing 225 short tons of ore through 
a pilot plant. The cobalt concentrate 
ranged in grade from 1. 5 to 3. 9 pcfe Co 
and recoveries ranged from 95 to 87 pet, 



respectively (.9)* Some research was 
conducted to sequentially float a Pb, Cu, 
and Co-Ni concentrate, but this technique 
could only be used with the high-grade 
portions of the ore. Benef iciation 
research for the Madison ore deposit 
appears to be complete. The technology 
has been tested on a pilot scale and no 
problems have been reported. 

The cobalt was extracted from the bulk 
sulfide concentrate with a sulfation 
roast followed by pressure leaching with 
H2SO4 (fig. 3). A continuous pressure 
leaching circuit was operated and 
extracted 96, 94, and 87 pet of the Co, 
Cu, and Ni. Minor amounts of Fe, Mn, and 
Zn impurities were also extracted. 
Nearly all of the lead remained in the 
leach residue. 

The copper in the pregnant solution was 
recovered by elect rowinning. The spent 
copper electrolyte was purified with 
ammonia and NaHS to remove the iron, 
residual copper, manganese, and zinc. 
Solvent extraction of the purified solu- 
tion using versatic acid recovered both 
cobalt and nickel. The organic was 
stripped into a chloride system, where 
another solvent extraction using tri n- 
octylamine (TNOA) separated the cobalt 
from the nickel. The cobalt was stripped 
from the organic and elect rowon, produc- 
ing cobalt metal and chlorine gas. The 
chlorine was reacted with hydrogen to 
produce hydrochloric acid, which was 
recycled back to the cobalt-nickel strip- 
ping operation. The nickel raffinate was 
treated to precipitate marketable nickel 
salts. This refining method is currently 
practiced by the Sumitomo Metals Mining 
Co. , LTD, Tokyo, Japan for the treatment 
of mixed sulfides (22). The operating 
costs for the Madison Mine are reported 
to be around $20/lb of Co produced. 

MISSOURI LEAD BELT DEPOSITS 

The typical Missouri lead belt ore 
contains 5.9 pet Pb, 1.1 pet Zn, 0.3 pet 
Cu, 0.02 pet Ni, and 0.015 pet Co, which 
is present as siegenite. These ores are 
a primary source of lead, zinc, and some 
copper. Based upon the annual production 
from Missouri lead belt mines, 2.5 
million lb of Co is mined, but not 



Ore 



H 2 S0 4 



Ammonia 
NaHS 







ir 










Flotation 








" 








Sulfation roast 








v 








Pressure leach 












i 


' 








Cu 
electrowinning 


k r* 


U 






p o 






i 


r 






— *. 


Purification 


M 


n, 


Zn 




* F 


Cu 












1 


r 










S* 










► Co + Ni 

1 








1 




» 








Stripping 




\\C\ A 




Waste 




MLI * 


H, 










1 










w 










r 


- 


SX 




Cobalt 
stripping 




Burner 
















k 




CI, 




j 


r 


^ 




,. J 






Precipitation 






Cobalt 
electrowinning 


i 


l 










Ni s 


alts 










\ 
C 


o 







FIGURE 3.— Madison Mine cobalt process. 



recovered. About 80 million lb of Co is 
present in the unmined ores. 

With benef iciation circuits already in 
place to recover lead, zinc, and some- 
times copper, research on the recovery of 
cobalt has been aimed at the mineral con- 
centrates and the tailings product. 
About 20 to 30 pet of the cobalt ends up 
in the chalcopyrite concentrate contain- 
ing 29 pet Cu, 0.9 pet Co, 1.2 pet Ni, 
6 pet Pb, and 0.5 pet Zn. The siegenite 
was finely interlocked with the 



chalcopyrite, which needed to be ground 
below 10-ym size to obtain liberation. 
After grinding, the chalcopyrite was 
refloated in several stages leaving the 
siegenite behind. The process was 
demonstrated on a 130-lb/h bleed stream 
from a lead mill copper circuit. About 
80 pet of the Co was recovered in a 
siegenite concentrate containing from 2.8 
to 3.8 pet Co and from 4 to 5 pet Ni. 

About half of the cobalt in the Misouri 
lead ores ends up in the flotation 



tailings, which contain 0.02 pet Co. 
Flotation with a mixture of amine and 
xanthate recovered 66 pet of this Co in 
a bulk sulfide concentrate containing 
0.11 pet Co. The process is currently 
being pilot tested at Cominco American, 
Inc's Magmont Mine, Iron County, 
Missouri. 

The remaining 20 to 30 pet of the 
cobalt is in the galena and sphalerite 
concentrates (23). Here the siegenite is 
finely interlocked with both the galena 
and sphalerite. Physical separation of 
the siegenite from these sources has not 
been tested. The benef iciation research 
for the Missouri lead ores appears to be 
nearing completion. About 20 pet of the 
ore's cobalt is recovered in the siegen- 
ite concentrate from the copper circuit. 
Bulk sulfide flotation has recovered 
33 pet of the ore's cobalt from the 
flotation tailings. However, this bulk 
sulfide concentrate is low grade. The 
potential for increasing this grade and 
maintaining cobalt recovery appears to be 
low, because the siegenite is interlocked 
with the other sulfide minerals at grain 
sizes below 5 ym (8_). 

Recovery of cobalt from the siegenite 
concentrate appears to be linked to 
marketing the concentrate to a cobalt 
refinery. The nickel content of this 
concentrate was lower than specifications 
for the world's Co-Ni-Cu refineries (22). 
However, it is possible that this concen- 
trate could be blended with higher grade 
nickel concentrates for cobalt extrac- 
tion. The next likely processing route 
is the method proposed by Anshutz Mining 
Corp., Fredricktown, MO, for the Madison 
Mine. This method should be tested on 
the siegenite concentrate to determine 
its feasibility and to obtain a prelimin- 
ary economic evaluation of the overall 
cobalt recovery process. 

No technology has been developed for 
the low grade, bulk sulfide concentrate 
from the lead mill tailings. The 
dolomite gangue will probably exclude 
acid leaching leaving ammoniacal leaching 
as the only potential extraction technol- 
ogy for this concentrate. Research needs 
to be conducted to determine the techni- 
cal feasibility of cobalt extraction from 
this concentrate. 



COPPER DEPOSITS 

Some of the primary copper ores of 
Utah, Arizona, and New Mexico contain 
cobalt at levels around 0.01 pet Co. A 
specific cobalt mineral has not been 
identified, but the cobalt appears to be 
more concentrated in the iron sulfide 
minerals. Although less than 10 pet of 
these deposits have significant 
concentrations of cobalt, with over 12 
billion st of ore in reserves, these 
deposits still could contain over 75 
million lb of Co (32). Possible by- 
product recovery of the cobalt has been 
identified in two portions of the copper 
production circuits. Pyrite and 
pyrrhotite in some flotation tailings 
contain 0.06 pet Co and spent copper 
leach solutions from some heap leaching 
operations contain 15 to 30 ppm Co. At 
least two deposits have been identified 
containing cobalt-bearing sulfides with a 
potential for producing 1.0 million lb of 
Co annually. Five leach solution sources 
have been identified with a potential 
for producing over 1.5 million lb of Co 
annually. 

Pyrite Concentrates 

Flotation of pyrite from the tailings 
of one Arizona mine produced a sulfide 
concentrate containing 0.06 pet Co. 
About 30 pet of the cobalt was recovered 
along with 50 pet of the copper. No 
research has been done to extract cobalt 
from this low-grade cobalt concentrate. 
Benef iciation and extraction methods need 
to be developed and preliminary economics 
of a process should be evaluated. The 
pyrite concentrates have nearly the same 
grade as most of the domestic laterites, 
but do not have the mining costs. 

Spent Copper Leach Solutions 

The second cobalt source from primary 
copper operations is the spent copper 
leach solutions. Heap leaching opera- 
tions extract the cobalt from the ore 
into a weak acid solution. The degree of 
extraction is not known. However, after 
recovery of the copper, the spent copper 
leach solution contained 15 to 30 ppm Co 



10 



(17). Ion exchange was used to recover 
the Co, Ni, and Cu from this solution 
(fig. 4). The loaded resin was sequen- 
tially stripped with (1) weak H 2 S0 4 to 
remove a portion of the Fe, Zn, and Al ; 
(2) 30 g/L H 2 S0 4 to recover the cobalt 
and nickel; and (3) NH 4 0H to remove the 
copper and residual nickel. Next, 
solvent extraction using di(2-ethylhexyl) 
phosphoric acid was used on the cobalt- 
nickel solution to remove the Fe, Zn, and 
Al impurities. Solvent extraction with 



CYANEX 272 was used on the cobalt-nickel 
bearing raffinate to recover the cobalt. 
The organic phase was stripped with spent 
cobalt electrolyte, followed by cobalt 
electrowinning. The process was demon- 
strated continuously with a multiple- 
compartment ion exchange unit and 
recovered 96 pet of the Co from the spent 

■'Reference to specific products does 
not imply endorsement by the Bureau of 
Mines. 



Spent Cu 
leach solution 



Ion exchange 



Weak H 2 S0 4 — ► 



Stripping 



H 2 S0 4 — ► 



-> Fe, Zn, Al 



Stripping 



NH4OH 



-► Co, Ni 



Stripping 



-►Cu, Ni 



SX 



Stripping 



-* Fe, Zn, Al 



SX 



Cobalt stripping 



Nickel 
precipitation 



Cobalt 
electrowinning 



NiC0 3 Co 

FIGURE 4.— Cobalt from spent Cu leach solution. 



11 



copper leach solution (18). An economic 
evaluation of the process from spent 
copper solution to cathode cobalt metal 
showed that, with credits for the copper 
and nickel values, cobalt could be 
produced for about $7.00/lb. 

The research on cobalt recovery from 
spent copper leach solutions is nearly 
complete. A long-term pilot program may 
be needed to establish resin life and the 
steady-state solution concentrations. An 
economic evaluation indicated that this 
cobalt source has the highest potential 
for domestic production. 

IRON DEPOSITS 

Cobalt is also found in some iron 
deposits, associated with sulfides. In 
the iron deposits of Cornwall, PA, 56 
million lb of Co is present at grades 
ranging from 0.02 to 0.056 pet Co. About 
40 to 60 pet of the ore is magnetite and 
3.5 pet pyrite (29). Historical process- 
ing of these ores recovered byproduct Cu, 
Co, Ag, and Au. In the 2 yr prior to 
closure of these iron mines, the cobalt 
grade dropped by 30 pet. However, it was 
not clear from that report whether the 
grade of cobalt in the sulfides dropped 
or if the sulfide content of the ore 
dropped. Before the iron mines closed, 
1.5 million lb of Co was produced 
annually. 

Benef iciation of these ores began with 
crushing and grinding followed by magnet- 
ic separation to recover a high-grade 
magnetite concentrate (fig. 5). The non- 
magnetic product was floated to recover a 
bulk sulfide concentrate, mostly pyrite, 
containing 0.7 to 1.4 pet Co. 

Cobalt extraction from the pyrite con- 
centrate began with roasting and shipping 
the calcine product to Pyrites Co. Inc. 
in Wilmington, DE. The calcine was per- 
colation leached with sulfuric acid for 
an average of 250 h. The Co, along with 
some of the Fe, Cu, and Mn, was dissolved 
into the solution. The extraction 
percentage was not reported. 



Ore 



Magnetic 
separation 



-+• Magnetite 



Flotation 
Tailings 



Pyrite 



Roast 



H 2 S0 4 



Na 2 C0 3 



Leach 



-► Fe 



Purification 



Na 2 C0 3 ► 

Cl 2 — ► 



-+■ Fe, Cu, Mn 



Precipitation 



_£ 



Calcined 

— J~~ 

Co 

oxide 



l-± 



Reduction 



T 

Co 



FIGURE 5.— Cobalt from Pennsylvania iron ore. 

The pregnant solution was purified with 
sodium carbonate, which precipitated the 
Fe, Cu, and Mn. After filtering, chlor- 
ine and more sodium carbonate were 
used to precipitate the cobalt. The 
filtered cobalt cake was either calcined 
to the oxide containing 70 pet Co or 
reduced to the Co metal with charcoal. 
The final metal powder contained 98 to 
99 pet Co. Over 99 pet of the cobalt in 
the solution was recovered (33). This 
process was used commercially up until 
1971 when the mining companies closed 
their iron mines, cutting off the supply 
of byproduct cobalt-bearing pyrite. If 
iron ore markets improve, this source of 
cobalt may resume production. 



12 



OXIDES 



LATERITE DEPOSITS 

The laterite deposits of northwestern 
California and southwestern Oregon are 
the result of extensive weathering of 
ultramafic serpentine rocks and contain 
over 96 million lb of Co. However, the 
deposits are located in a wilderness 
preserve. The average grade is 0.08 pet 
Co, but varies throughout the deposits 
from 0.06 to 0.25 pet Co. There is also 
0.5 to 1.2 pet Ni, 2 pet Cr, and 0.3 pet 
Mn in these deposits. Nearly all of the 
Co is found in the manganese mineral, 
which makes up only 1 pet of the ore. 
Well over a third of the ore is goethite, 
which contains most of the nickel. The 
remainder of the ore is quartz, hematite, 
magnetite, and chromite (5). 

Ore benef iciation of the laterite 
deposits has been limited to crushing and 
screening out some of the coarse pieces 
of quartz (30). The grain size of the 
goethite and manganese mineral is very 
small, which would require fine grinding 
of the ore to obtain liberation. No 
successful benef iciation techniques have 
been developed for this fine-grain 
mineral separation. Even with a perfect 
separation, the Co and Ni grades would 
only be 2.5 times larger than in the ore. 
Considering the limited benef iciation 
potential and the lack of technology, 
these laterite ores will not be signifi- 
cantly beneficiated before the extraction 
process. 

The extraction of cobalt from laterites 
has received considerable research atten- 
tion. The two major techniques for 
cobalt extraction are reduction roast- 
ammoniacal leach (fig. 6) and H2SO4 leach 
(fig. 7). The basic reduction roast- 
ammoniacal leach process shown in figure 
6 began with drying and crushing the ore 
to minus 2-cm size. Selective reduction 
of the nickel and cobalt with H2 and CO 
was conducted in multiple hearth furnaces 
at 700° to 760° C. The reduced ore was 
leached with an aerated solution of 
ammonium hydroxide and ammonium carbon- 
ate. The nickel and cobalt were leached 
out while the iron remained insoluble. 
This process, commonly called the Caron 



process, was used in Australia and the 
Philippines. Only about 50 pet of the 
cobalt and 75 to 80 pet of the nickel 
were recovered. 

The Bureau's reduction roast-ammoniacal 
leach process was tested on laterite ores 
(30). Crushed pyrite was added to the 
ore followed by multiple-hearth roasting 
at 525° C with pure CO gas. The nickel 
and iron combined to form ferronickel and 
the copper and cobalt were reduced to the 
individual metals. After cooling, the 
calcine was finely ground and leached 
with ammonium hydroxide, ammonium sul- 
fate, and oxygen to dissolve Co and Ni as 
ammine complexes. The process was 
demonstrated in a 500 lb/d integrated, 
continuous operation using an ore sample 
containing 0.2 pet Co and 0.97 pet Ni. 
About 85 pet of the Co and 90 pet of the 
Ni were recovered in the pregnant leach 
solution. However, with ore samples 
containing 0.05 to 0.1 pet Co (average, 
0.08 pet), a 4. 5-st/d pilot plant test of 
this process only extracted 67 pet of the 
Co and 83 pet of the Ni (3_1> 

Laterites were also processed with a 
H2SO4 leach process (fig. 7) (_3). The 
ore containing 0. 08 pet Co and 1 pet Ni 
was slurried and pumped to leaching 
towers where it was leached with H2SO4 at 
200° to 250° C under more than 500-psi 
pressure. The process dissolved cobalt, 
nickel, and magnesium, while the iron was 
hydrolyzed and precipitated. In labora- 
tory tests, about 85 to 90 pet of the 
cobalt and 90 to 95 pet of the nickel was 
extracted. California Nickel Corp. , San 
Francisco, CA completed pilot-plant work 
on this process and AMAX proposed to 
employ this process for the New Caledonia 
C0FREMI laterite operations to recover 
nickel and cobalt (2^). To minimize acid 
consumption, the process is generally 
restricted to low-magnesia ores (less 
than 5 pet MgO). The California and 
Oregon laterites contain 4. 3 to 7. 5 pet 
MgO as serpentine and chlorite. 

The solution refining process depends 
upon the extraction method. For the 
ammoniacal solutions, NH4H2PO4 was added 
to precipitate Mg and Mn as (Mg, Mn) 
NH4PO4, a fertilizer byproduct. After 



Laterite 



Crush 



13 



FeS 3 



NH 4 H 2 P0 4 



Dry 



Calcine 
- 525° C 
reduction 



Cool 



Grind 



Leach 



Liquid - solids 
separation 



Product liquor 



Impurity removal 



CO (fuel) 



Coke 



C0 2 
CO 



CO producer 



Ash 
Makeup NH 4 OH, 
(NH 4 )2S0 4 , H 2 



2 



Leach solution 

NH 3 L (NH 4 )2S0 4 

solution 



Recovery NH 3 

and (NH 4 hS0 4 

solution 



Solution 



H,0 



Tailings 
wash and filter 



NH 4 (Mg, Mn)P0 4 



NH 3 stripping 



H 2 



Gangue 



NH, 



Ni-Cu solvent 
extraction 



Spent 

electrolyte 

recycles 



Absorber 



Raffinate 



Co solvent 
extraction 



Electrolytes 

+ i i 



Electrowinning 



Spent 

electro- 



11 I ■ eiectro- 

I I iyte 

▼ ▼ recycle ▼ 



Raffinate recycle 



Ni Cu Co ZnS0 4 

FIGURE 6.— Reduction roast-ammoniacal leach process for laterites. 



14 



H 2 S0 4 ► 



Ore 



Pressure leach 



H 2 S 



Purification 



Mn 
concentrate 



Acid <- 



Co+Ni 
precipitation 



Mn 
precipitation 



MgS0 4 H 2 
crystallization 



MgO recovery 
MgO 



-> Zn, Cu 



Co+Ni 



Ammonia leach 



Ammonia 



± L 



SX 



Stripping 



Ni 
electrowinning 



Ammonia 
stripping 



Ni 



Co leach 






Co 
electrowinning 



T 

Co 



FIGURE 7.— Sulfuric acid leach process for laterites. 



stripping excess ammonia from the solu- 
tion with steam, solvent extraction was 
used to remove both the Ni and the Cu. 
The nickel was stripped from the organic 
with spent nickel electrolyte followed by 
a electrowinning the nickel. The copper 
was then stripped from the organic with 
spent copper electrolyte followed by 
copper electrowinning. The Ni-Cu raffin- 
ate was neutralized precipitating the 
residual Cu, Mg, Zn, Mn, and Ni. The 
cobalt in the solution was reduced from 
Co 3+ to Co 2+ with cobalt shot and solvent 
extraction used to recover the cobalt. 
The organic was stripped 
cobalt electrolyte and the 
electrowon. This process 
strated in a pilot plant, 
metal only 65 pet of the 



with spent 
cobalt was 
was demon- 
From ore to 
cobalt and 



83 pet of the nickel were recovered in 
the cathode products (31 ). 

For the H2SO4 extraction refining 
process, H 2 S was added to precipitate Cu 
and Zn as sulfides. Next the Ni and Co 
were precipitated with lime. Then the 
Ni-Co precipitate was leached with 
ammonia solution, followed by solvent 
extraction. The organic phase was 
stripped with spent nickel electrolyte 
and the nickel was electrowon. The 
raffinate was treated with steam to strip 
the ammonia and then the cobalt was 
precipitated. The cobalt precipitate was 
leached with spent cobalt electrolyte 
followed by cobalt electrowinning. A 
5-st/d pilot plant was operated to demon- 
strate this technique. From ore to metal 
over 90 pet of the Co and Ni were 



15 



recovered (2^). Research on laterite 
processing appears to be complete, how- 
ever, the economics are not favorable. 
The residues have been evaluated and 
proved to be nontoxic to the environment 
(25), but development of these deposits 
will face significant opposition from 
environmentalists. 

MANGANESE SEA NODULES 

Manganese nodules are found over large 
areas of the ocean floor at depths of 
around 2,000 to 18,000 ft. They range 
from 0.25 to 3 in. in size and contain 25 
to 30 pet Mn, 1.0 to 1.5 pet Ni, 0.5 to 
1.0 pet Cu, and 0.25 pet Co (26). Large 
concentrations are found in the east 
central Pacific area. Only a small por- 
tion of nodule samples were taken from 
the U.S. Exclusive Economic Zone (EEZ ) 
located off the Hawaiian shore and other 
islands of the Trust Territory of the 
Pacific Islands and affiliated terri- 
tories. Estimates as high as 1 billion 
lb of Co have been made for this 
resource. 

The mineralogy is basically fine 
grained oxides mixed with layers of 
silicious gangue minerals (13). Although 
a number of problems remain to be solved, 
the mining technology proposed for deep- 
sea nodules has been a hydraulic air 
suction system or a continuous-line 
bucket system. Much of the prototype 
equipment has already been designed, 
patented, built, and tested (15). How- 
ever, international politics and many 
legal problems appear to have stalled the 
development of this resource. 

The fine grain mineral structure of the 
nodules precludes any effective benefici- 
ation except for physically separating 
the nodules from the bottom sediment by 
screening 

Extraction Methods 

A great deal of research has been 
conducted to develop methods for extract- 
ing the Mn, Cu, Ni, and Co from the 
nodules (12). Five of the most promising 
techniques for cobalt extraction are 
(1) gas reduction and ammoniacal leach, 



(2) cuprion ammoniacal leach, (3) high- 
temperature and high-pressure H 2 S0 4 
leach, (4) reduction and HC1 leach, and 
(5) smelting and H2SO4 leach. 

Gas Reduction and Ammoniacal Leach 
Process 

First, the nodules were ground to 65- 
mesh size and dried followed by high- 
temperature (625° C) reduction of mangan- 
ese dioxide to manganese oxide by CO gas 
(fig. 8). After cooling to 40° C, the 
Cu, Ni, and Co were dissolved with an 
oxidizing ammoniacal ammonium carbonate 
leach (Caron process). About 90 pet of 
the Cu, 90 pet of the Ni, and 70 pet of 
the Co were dissolved into the solution. 

Cuprion Ammoniacal Leach Process 

The nodules were ground to 65-mesh size 
and reduced with cuprous ions (Cu + ) in a 
ammoniacal solution at 50° C (fig. 9). 
The manganese dioxide was reduced to 
manganese oxide. The cobalt was leached 
with a strong solution of ammonia and 
carbon monoxide. Next, the slurry was 
oxidized with air to oxidize the soluble 
ions and precipitate the iron. Only 
50 pet of the Co was recovered, but 
90 pet of the Cu and Ni was recovered 
(1_). This extraction technique is not 
appropriate for cobalt recovery. 

High-Temperature, High-Pressure H2SO4 
Leach Process 

The nodules were ground to 65-mesh size 
and leached with H 2 S0 4 at 245° C and 
500-psi pressure (fig. 10). The recovery 
of Co, Cu, and Ni was 90, 95, and 95 pet, 
respectively. Small amounts of Fe, Mn, 
and Zn were also dissolved. 

Reduction and HC1 Leach Process 

The nodules were ground to 65-mesh size 
and dried, followed by high-temperature 
(500° C) gaseous HC1 reduction of mangan- 
ese dioxide to manganese chloride (fig. 
11). This reaction also produced chlor- 
ine gas that reacted with the other metal 
oxides, forming metal chloride salts. 



16 







Ni 

t 








Cu 

t 


Co 

▲ 




Nodules 

1 


Carbon monoxide 

i 


i-> 


Electro- 
winning 


•4- 


r* 


Electro- 
winning 










Grinding 




Chemical 
reduction 
































Ni 
stripping 




Cu 
stripping 


«- 






i 


i 


IS«« 


V 








SolUuun 








recycle 


Leaching - 

reduced pulp 

preparation 




















































Ni- 
ion 


Cu liquid 
exchange 






Co 

recovery 






i 














' 














Waste 
containment 




Leaching - 
liquid-solid 
separation 


Metal 
so 


► 


w 




- bearing 
ution 






1 






i 


k 




1 




L 






■ •*. 






Ammonia 
recovery 








Makeup 






water 


T 

Mak 
amm 


i 

eu 
on 


P 
a 


i 































FIGURE 8.— Gas reduction and ammoniacal leach process for sea nodules. 









Ni 

t 








Cu 

t 


Co 

▲ 




Nodules 




co 2 -co 




-► 


Electro- 
winning 


■4- 


r> 


Electro- 
winning 










Grinding 
and drying 




Chemical 
reduction 


























-► 


Ni 
stripping 


- 


Cu 
stripping 








i 


L 




i 


' 














Leach 




















































Ni- 
ion 


Cu liquid 
exchange 






Co 
recovery 






' 


r 




















Waste 
containment 




Liquid - solid 
separation 


■ w 


w 




Metal-bearing 
solution 








T 

H,S 






i 


i 




j 


. 










Ammonia 
recovery 








Makeup 










water 




T" 

Mak 
amm 


i 

eu 
on 


P 
a 


l 





























FIGURE 9.— Cuprion ammoniacal leach process for sea nodules. 



FIGURE 10.— High-temperature, high-pressure H 2 S0 4 leach process for sea nodules. 



Grinding 

and 
drying 



Nodules 



Hydrochloric 
acid 



Hydro- 
chlorination 
- reduction 



H 2 



Hydrochloric 

acid-chlorine 

recovery 



Hydrolysis 



Water 
storage 



Cu 
liquid ion 
exchange 



Leach 

liquid - solid 

separation 



Waste 
containment 



Electro- 
winning 



Co 
liquid ion 
exchange 



H 2 S 
_i_ 



Co 
recovery 



Ni 
liquid ion 
exchange 



Electro- 
winning 



Evaporation - 
crystalization 



Fused 

salt 

electrolysis 



FIGURE 11.— Reduction and HC1 leach process for sea nodules. 



17 











Cu 


Ni 


Grinding 


4 


Nodules 

k 


t 






t 






Electro- 
winning 


h. 


Electro- 
winning 






f 




H 2 














J 
















Leaching 


4 




H2SO4 






Cu 
liquid ion 
exchange 












Ni 
liquid ion 
exchange 




PH 
adjustment 




^ — l 


Neutralization 


fe 


^ — ' 










t 


»- 


W 










i 




















Liquid - solid 
separation 










T 








i 


i 




















1 ■ 


i 


















1 






Ammonia 
recovery 




Co 
recovery 


4 


Y 








Waste 
containment 








k n n 






^ 




4 Lime 


~ \*o 






t t 




t 














Makeup Makeup H 2 S 
ammonia water 





-> Cu 



Co 



-►Mn 



Water was sprayed to cool the product and 
caused the iron to form insoluble ferric 
hydroxide. After cooling, the mixture 
was leached with HC1 acid. Over 99 pet 



of the Co was extracted along with 96, 
99, and 94 pet of the Cu, Ni, and Mn, 
respectively. 



18 



Smelting and H 2 S0 4 Leach Process 



Refining Process 



The nodules were dried, roasted with 
coke and CO gas at 725° C, and smelted 
with silica flux at 1,425° C (fig. 12). 
The slag, containing Mn, Fe, and Si0 2 , 
was removed and subsequently used for 
f erromanganese production. The manganese 
alloy containing the Fe, Cu, Ni, and Co 
was converted by adding sulfur and heat- 
ing to form a matte and an iron alloy. 
More silica flux was added and the matte- 
iron alloy was blown with air to lower 
the iron content of the matte to 5 pet. 
The finished matte was granulated, ground 
to 325-mesh size, and pressure leached 
with H2SO4 at 150 psi and 110° C. About 
90 pet of the Co was dissolved in the 
solution along with 90 pet of the Cu and 
Ni. The Co recovery was uncommonly high 
for a matte smelting process. 



Refining of the leach solutions depends 
upon the solution type such as ammonia- 
cal, acid chloride, or acid sulfate. 

Ammoniacal Solution Refining 

For the gas reduction and ammoniacal 
leach process and cuprion process, the 
pregnant liquor passed through a solvent 
extraction circuit where the copper, 
nickel, and some ammonia were removed. 
Most of the ammonia was removed from the 
organic phase by washing with a weak 
aqueous ammonia solution. The organic 
phase was cleansed of the residual 
ammonia with a slightly acidic ammonium 
sulfate solution. The nickel was care- 
fully stripped from the organic phase 
using spent nickel electrolyte containing 



Reducing gas 



Chemical 




Grindii 


■tg and 


< Nodules 




reduction 




drying 




1 


' 








Electric 
furnace 
smelting 


Ferro- 

manganese 

reduction 




Ferromanganese 










3 


Gyr. 

r i 


»sum 
2 

> 1 


3 


r 




Cu 

t 




Ni 

t 




Oxidizing 
sulfiding 


Waste 
treatment 




Electro- 
winning 




Electro- 
winning 








•w 






i — w 


«— ' 




























r 




pH 
adjustment 






Cu 
liquid ion 
exchange 






Neutralization 






Ni 
liquid ion 
exchange 


1 


1 










Leaching- 

liquid - solid 

separation 










4- 




















z 










1 






















r 


1 




Ammonia 
recovery 




Co 
recovery 


^ 












Waste 


-* 


CO 


ntai 


nment 










T 





Co 



H 2 S 



FIGURE 12.— Smelting and H 2 S0 4 leach process for sea nodules. 



19 



40 g/L H 2 S0 4 . Then the nickel was 
electrowon from the nickel-rich electro- 
lyte. The copper was stripped from the 
nickel-free organic phase using a spent 
copper electrolyte containing 160 g/L 
H 2 S0 4 . The copper was electrowon from 
the copper-rich electrolyte and the 
metal-free organic was recycled to the 
solvent extraction circuit. 

The LIX raffinate was treated with 
ammonium hydrosulf ide, which precipitated 
the Co along with residual Cu, Ni, and 
Zn. The sulfide mixture was pressure 
leached with air to preferentially 
dissolve the Ni and Co, leaving the Cu 
and Zn sulfides behind in the residue. 
After filtering off the residue, hydrogen 
sulfide was used to remove the residual 
Zn and Cu impurities. Then the solution 
was heated in an autoclave with hydrogen 
to reduce the Ni. The nickel powder was 
washed, dried, and briquetted. The 
cobalt sulfate solution was concentrated 
in an evaporator-crystallizer to precipi- 
tate the Co and residual Ni with ammonium 
sulfate. The salts were redissolved with 
a strong ammonia solution and oxidized 
with air, converting the Co 2+ to Co 3+ . 
The remaining nickel was precipitated 
with acid and removed from the solution. 
The nickel-free solution was reduced with 
hydrogen in an autoclave, producing 
cobalt powder, which was dried and 
briquetted. About 98 pet of the cobalt 
was recovered in the refining process. 
The overall cobalt recovery from nodule 
to cobalt powder was 89 pet. 

Acid Sulfate Refining 

The pregnant solution from the high- 
temperature, high-pressure H 2 S0 4 leach 
process and the smelting and H2SO4 leach 
process were refined by the same method. 
First, the excess acid was neutralized 
with limestone to lower the H2SO4 content 
to 0.5 g/L. After filtering off the 
precipitated gypsum, the solution passed 
through solvent extraction with LIX to 
transfer the copper to the organic phase. 
The copper was stripped with spent copper 



electrolyte containing 160 g/L H 2 S0 4 . 
Then the copper was electrowon from the 
copper electrolyte. The copper raffinate 
was adjusted with ammonia to a pH of 4. 
During neutralization the tank was also 
aerated, causing the Fe, Mn, Mg, and Al 
impurities to precipitate. After centri- 
fuging out the solids, solvent extraction 
collected the nickel in the organic 
phase. The nickel was stripped from the 
organic phase with spent nickel electro- 
lyte. Then the nickel was electrowon 
from the nickel-rich electrolyte. The 
nickel raffinate was purified by the 
method described earlier for the LIX 
raffinate in the ammoniacal refining 
method. The final product was briquetted 
cobalt metal powder. Again, 89 pet of 
the cobalt was recovered from the 
nodules. 

Acid Chloride Refining 

The pregnant liquor was passed through 
a solvent extraction circuit to remove 
the Cu. The organic phase was stripped 
with spent copper electrolyte containing 
160 g/L H 2 S0 4 . Then the copper was 
electrowon from the copper-rich electro- 
lyte. The pH of the raffinate from the 
copper extraction was adjusted to 4 with 
NaOH. The cobalt and some of the mangan- 
ese were recovered by solvent extraction 
with TNOA. Both the cobalt and manganese 
were stripped from the organic and then 
the cobalt was precipitated with H 2 S. 
This precipitate was treated by the same 
method reported earlier for the ammonia- 
cal refining method, beginning with the 
pressure leaching of the mixed sulfide 
precipitates. The cobalt was recovered 
as a briquetted metal powder. About 
99 pet of the cobalt was recovered from 
the nodules. The cobalt-free manganese 
solution went to the manganese recovery 
circuit. The cobalt-free raffinate was 
treated by solvent extraction to recover 
the nickel. The nickel was stripped from 
the organic phase with spent nickel 
electrolyte and recovered by electro- 
winning. The nickel-free raffinate was 



20 



combined with the cobalt-free raffinate 
and treated with H 2 S to precipitate any 
residual Co, Ni , or Cu as sulfides. The 
remaining solution was evaporated, 
crystallizing manganese chloride salt. 
The salt was dried and fed to a high- 
temperature (1,300° C) f used-salt elec- 
trolysis furnace where molten manganese 
metal was formed and periodically tapped. 
About 96, 99, and 94 pet of the Cu, Ni, 
and Mn were recovered from the nodules. 

Research on manganese sea nodule 
processing is complete. The residues 
were environmentally acceptable and the 
process waste waters blended safely with 
seawater (9_). At least five process 
options are available, but the economics 
are not favorable. At a price of $17/lb, 
the calculated rate of return on invest- 
ment was 4 to 6 pet. However, at least a 
30-pct rate of return on investment is 
required to attract the necessary venture 
capital for sea nodule production. 

MANGANESE SEA CRUSTS 

Cobalt is also found in offshore manga- 
nese crust deposits located in the EEZ of 
Hawaii and Trust Territory of the Pacific 
Islands and affiliated territories. 
Cobalt-rich deposits on seamounts and 



slopes have been estimated to contain up 
to 25 billion lb of Co. The average 
cobalt content of these crusts is 1.0 pet 
and they also contain 0.5 pet Ni, and 15 
to 25 pet Mn ((>)• These deposits, 1-in 
thick layers, present a problem for 
selective mining. Equipment is being 
developed to break up the crust layer and 
leave most of the substrate behind, but 
this technology is still being tested. 
The mineralogy of the Pacific sea crusts 
is very similar to the Pacific nodules. 
Some places contain both crust and 
nodules. Even with a good selective 
mining system, benef iciation will be 
needed to separate the crust minerals 
from the substrate. Preliminary flota- 
tion tests have recovered 92 pet of the 
cobalt in a crust mineral concentrate. 
This research is not complete. 

The extraction technology is expected 
to be identical to the sea nodule extrac- 
tion methods. This resource still needs 
exploration, because very few samples 
have been taken from U.S. EEZ. The 
mining and benef iciation technology needs 
to be developed. The extraction technol- 
ogy is well developed, but the overall 
costs from mining through cobalt process- 
ing will be high. 



DISCUSSION 



The full potential for cobalt produc- 
tion from domestic sources is at least 
19.4 million lb of Co per year. With the 
offshore manganese deposits, the amount 
doubles. A summary of each deposit's 
characteristics, benef iciation, extrac- 
tion, and refining results is presented 
in table 1. Also, included is a six- 
level technological assessment of the 
research and development status for the 
benef iciation and cobalt processing of 
each ore. 

The technology for Blackbird, Madison 
Mine, iron ore pyrite, laterites, and 
manganese sea nodules is nearly complete- 
ly developed, but the economics are not 
favorable. Small technological improve- 
ments will have virtually no effect upon 
the economics of these operations. The 
most promising source of cobalt is the 



spent copper leach solution. With all 
the mining and benef iciating costs paid 
by the primary copper production, cobalt 
can be extracted at a relatively low 
cost. The Missouri lead belt ores are 
also promising. Siegenite benef iciation 
from the copper concentrate is well 
developed, however the technology for 
extracting and refining the cobalt from 
the siegenite concentrate has not been 
fully explored. A secure domestic 
processing plant is needed to produce 
cobalt from the siegenite concentrate. 

It appears that benef iciation can not 
increase the bulk sulfide grade from the 
lead mill tailings without losing 
recovery. Because of the dolomite 
gangue, cobalt processing technology 
for this concentrate will probably be 
limited to an ammoniacal leach. The 



21 



applicability of this technology needs to 
be established so that the technical 
feasibility and economics of cobalt 
recovery from this source can be 
evaluated. 

The pyrite from some primary copper 
operations also shows potential for sig- 
nificant cobalt recovery. The technology 
for benef iciation and cobalt processing 
still needs research, but without the 
mining cost burden, this cobalt source 
should be one step closer to favorable 
economic production of cobalt. 

Both the Duluth Gabbro and Yakobi 
Island deposits show potential for 
cobalt. The benef iciation technology for 
the Duluth Gabbro deposit is complete, 
but the cobalt processing technology 
needs to be established on the Co-Ni 
concentrate. Cobalt production from this 
deposit appears to be many years into the 
future. For each pound of Co produced 
from this deposit, 95 lb of Cu and 14 lb 
of Ni are also produced. The present 
market conditions for copper offer no 
incentive for exploiting this deposit for 
copper. 

The deposit on Yakobi Island is in a 
similar predicament. The benef iciation 
and processing technologies need to be 
explored. With 6 lb of Cu and 10 lb of 
Ni produced for each pound of cobalt, 
there is also little hope for development 
in the near future. 



Economic evaluation of cobalt deposits 
should help to focus research on the 
deposits with the best future for econom- 
ical production. A Bureau cobalt avail- 
ability study on cobalt from world 
resources indicated that the mining and 
benef iciation costs are a substantial 
portion of the operating costs (20). 
That is why byproduct recovery from 
present copper, lead, and zinc opera- 
tions, where the mining and benef iciation 
costs are already paid, have the highest 
potential for immediate cobalt produc- 
tion. Over 5 million lb of Co per year 
could be produced from these operations. 
Research on these sources should be given 
the highest priority. Even with virtual- 
ly unlimited offshore cobalt resources, 
the high mining and ore transportation 
costs associated with the offshore mining 
will postpone the development of these 
resources for many years (14). 

Under the present economic conditions 
cobalt could be made available within 2 
to 3 yr only as a byproduct from present 
operations. With higher but stable 
commodity prices, many of the other 
cobalt resources could be available in 4 
to 5 yr. In the event of a national 
emergency, the present cobalt stockpile 
could meet U.S. cobalt demand for 3 yr 
(4_) or with restrictions the stockpile 
could meet essential U.S. needs for up to 
6 yr (22). 



22 



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CONCLUSIONS 



23 



Byproduct recovery of cobalt from 
present copper, lead, and zinc operations 
has the highest potential for immediate 
cobalt production. This includes spent 
copper leach solutions from primary 
copper production, the siegenite concen- 
trate from Missouri lead-zinc-copper 
production, and pyrite concentrates from 
some primary copper mines. 

Technology research for the Blackbird, 
Madison Mine, iron ore pyrite, laterites, 
and manganese sea nodules is nearly 
complete. Development depends upon 
favorable economic conditions. Addi- 
tional research on these resources should 
be limited to approaches that promise to 
cut the total processing costs by over 
50 pet. 

Research on the Duluth Gabbro and 
Yakobi Island deposits is not complete 



and they will not be developed until 
favorable copper and nickel markets 
reappear. This will be many years in the 
future, so that research is very long 
range. 

Except for benef iciation of the sea 
crusts, research on sea nodules and sea 
crusts is complete. Development of these 
resources is many years off due to 
economic, legal, and political problems. 

Solution mining or heap leaching should 
be explored for the lead mill tailings, 
laterites, and Duluth Gabbro deposits as 
an alternative lower cost mining and 
extraction process. Once the cobalt is 
in the solution, the Bureau's spent 
copper leach solution refining technology 
could be used to recover the cobalt. 



REFERENCES 



1. Agarwal, J. C, H. E. Barner, 
N. Beecher, D. S. Davies, and R. N. Kust. 
Kennecott Process for Recovery of Copper, 
Nickel, Cobalt, and Molybdenum From Ocean 
Nodules. Pres. at Annual Meeting, AIME, 
Denver, CO, Feb-28-Mar. 2, 1978, Preprint 
78-B-89, 6 pp. 

2. American Metal Market (New York). 
AMAX, French Partner Offering Leach 
Process. V. 93, No. 51, Mar. 15, 1985, 
p. 7. 

3. . New Way of Extracting 

Cobalt, Manganese Found. V. 93, No. 66, 
Apr. 5, 1985, p. 7. 

4. American Society for Metals. 
Quality Assessment of National Defense 
Stockpile Cobalt Inventory. ASM, 1983, 
40 pp. 

5. Chandra, D. , C. 0. Ruud, and 
R. E. Siemens. Characterization of 
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and X-Ray Techniques. BuMines RI 8835, 
1983, 12 pp. 

6. Clark, A. L. , P. Humphrey, 
C. J. Johnson, and D. K. Pak. Cobalt - 
Rich Manganese Crust Potential, EEZ: 
U.S. Trust and Affiliated Territories in 
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and International Minerals, 1985. 



7. Clifford, R. K. , and L. W. Higley, 
Jr. Cobalt and Nickel Recovery From 
Missouri Lead Belt Chalcopyrite Concen- 
trates. BuMines RI 8321, 1978, 14 pp. 

8. Cornell, W. L. , D. C. Roltgrefe, 
and F. H. Sharp. Recovery of Cobalt and 
Other Metal Values From Missouri Lead Ore 
Concentrator Tailings. Paper in Recycle 
and Secondary Recovery of Metals, ed. by 
P. R. Taylor, H. Y. Song, and N. Jarrett. 
Metall. Soc. AIME, Warrendale, PA, 1985, 
pp. 675-682. 

9. DeCuyper, J. Concentration of 
Cobalt Ores, in International Conference 
on Cobalt. Brussels: Benelux Metallur- 
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10. Harris, G. B. , S. Monette, and 
R. W. Stanley. Hydrometallurgical Treat- 
ment of Blackbird Cobalt Concentrate. 
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AIME, 1983, pp. 139-150. 

11. Haynes, B. W. , S. L. Law, 
D. C. Barron, G. W. Kramer, R. Maeda, and 
M. J. Magyar. Pacific Manganese Nodules: 
Characterization and Processing. BuMines 
B 679, 1985, 44 pp. 



60142 1S9 



24 



12. Haynes, B. W. , S. L. Law, and 
R. Maeda. Updated Process Flowsheets for 
Manganese Nodule Processing. BuMines 
IC 8924, 1983, 100 pp. 

13. Haynes, B. W. , S. L. Law, and 
D. C. Barron. Mineralogical and Elemen- 
tal Description of Pacific Manganese 
Nodules. BuMines IC 8906, 1982, 60 pp. 

14. Hillman, C. T. , and B. B. Gosling. 
Mining Deep Ocean Manganese Nodules. 
Description and Economic Analysis of 
Potential Venture. BuMines IC 9015, 
1985, 19 pp. 

15. Hillman, C. T. Manganese Nodule 
Resources of Three Areas in the Northeast 
Pacific Ocean: With Proposed Mining- 
Benef iciation Systems and Costs. A 
Minerals Availability System Appraisal. 
BuMines IC 8933, 1983, 60 pp. 

16. Jansons, U. Cobalt Content in 
Samples From the Omar Copper Prospect, 
Baird Mountains, Alaska. BuMines Mineral 
Land Assessment 109-82, 1982, 16 pp.; 
available upon request from the Inter- 
mountain Field Operations Center, Denver, 
CO. 

17. Jeffers, T. H. Separation and 
Recovery of Cobalt From Copper Leach 
Solutions. J. Met., v. 11, No. 1, Jan. 
1985, pp. 47-50. 

18. Jeffers, T. H. , K. S. Gritton, and 
P. G. Bennett. Recovery of Cobalt From 
Spent Copper Leach Solution Using Contin- 
uous Ion Exchange. Paper in Recycle and 
Secondary Recovery of Metals, ed. by 
Patrick R. Taylor, Hong Y. Song, and 
Noel Jarrett. Metall. Soc AIME, Warren- 
dale, PA, 1985, pp. 609-621. 

19. Kirk, W. S. Cobalt. Ch. in 
BuMines Minerals Yearbook. 1985, v. 1 
pp. 295-304. 

20. Mishra, C. P., C. D. Sheng-Fogg, 
R. G. Christiansen, J. F. Lemons, Jr., 
and D. L. De Giacomo. Cobalt Availabil- 
ity-Market Economy Countries. A Minerals 
Availability Program Appraisal. BuMines 
IC 9012, 1985, 33 pp. 

21. Mobbs, D. C. , and D. M. Mounsey. 
Use of Caro's Acid in the Separation of 
Cobalt and Nickel. Trans. Inst. Min. 
Metall., Sec. C: Miner. Process Extr. 
Metall., v. 90, Sept. 1981, pp. C103- 
C110. 

22. National Materials Advisory Board. 
Cobalt Conservation Through Technological 



Alternatives. Natl. Acad. Sci. , NMAB- 
406, 1983, 115 pp. 

23. Paulson, D. L. , W. M. Dressel, and 
R. M. Doerr. Cobalt and Nickel Recovery 
From Missouri Lead Ores. Proceedings of 
the Seventh Mineral Waste Utilization 
Symposium, October 20, 1980. BuMines and 
IIT Res. Inst. Chicago, IL, 1980. 

24. Peterson, G. R. , D. I. Bleiwas, 
and P. R. Thomas. Cobalt Availability — 
Domestic. A Minerals Availability System 
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31 pp. 

25. Powers, L. A., and R. E. Siemens. 
Examination of Effluents Generated From 
Processing Domestic Laterites. BuMines 
RI 8797, 1983, 13 pp. 

26. Rowland, R. W. , M. R. Goud, and 
B. A. McGregor. The U. S. Exclusive 
Economic Zone - A Summary of Its Geology, 
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U.S. Geol. Surv. Circ. 912, 1983, 
29 pp. 

27. Schluter, R. B. , and W. M. Mahan. 
Flotation Responses of Two Duluth Complex 
Copper-Nickel Ores. BuMines RI 8509, 
1981, 24 pp. 

28. Shah, I. D. , P. L. Ruzzi, and 
R. B. Schluter. Low-Iron Cu-Ni-Co Matte 
From Duluth Complex Sulfide Concentrate 
by Direct Smelting. BuMines RI 8752, 
1983, 10 pp. 

29. Sibley, S. F. Cobalt. Ch. in 
Mineral Facts and Problems, 1980 Edition. 
BuMines B 671, 1980, pp. 199-214. 

30. Siemen, R. E. , and J. D. Corrick. 
Process for Recovery of Nickel, Cobalt, 
and Copper From Domestic Laterites. Min. 
Congr. J. , Jan. 1977, pp. 29-34. 

31. U0P, Inc. , Mineral Sciences Div. 
New Procedure for Recovering Nickel and 
Cobalt From Western Laterites. Economic 
Feasibility (contract J0285021). BuMines 
OFR 68-82, 1982, 169 pp.; NTIS PB 82- 
245945. 

32. Vhay, J. S., D. A. Brobst, and 
A. V. Heyl. Cobalt Ch. in United States 
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Prof. Paper 820, 1973, pp. 143-155. 

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and Problems, 1965 Edition. BuMines 
B 630, 1965, pp. 241-247. 



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